Process for recovering metal values from jarosite solids

ABSTRACT

This invention concerns a process for recovering metal values from jarosite-containing materials by leaching with a calcium chloride solution at a temperature above about the atmospheric boiling point of the solution and under at least the autogenous pressure.

FIELD OF THE INVENTION

This invention relates to a process for recovering metal values from thejarosite containing wastes from electrolytic zinc recovery plants.

BACKGROUND OF THE INVENTION

Electrolytic zinc processes are used to treat complex zinc-containingores that cannot readily be treated by pyrometallurgical recovery. Theusual steps in such an electrolytic process include: (a) concentratingthe zinc ores; (b) roasting the zinc concentrate to eliminate sulfur andproduce zinc calcine; (c) leaching the zinc calcine to provide an impurezinc sulfate solution; (d) separating the iron present usually byforming a jarosite precipitate; (e) purifying the zinc sulfate solution;and (f) subjecting the zinc sulfate solution to electrolysis to recoverthe zinc metal. Such a process is described in U.S. Pat. No. 4,128,617(1978) of DeGuire et al. which is incorporated herein by reference. Asimplified process flow sheet for an electrolytic zinc plant is shown inFIG. 1.

Additional details of various process modifications can be found in "TheEncyclopedia of Chemical Technology", Kirk-Othmer, Vol. 24, pp. 812-824,3rd Ed, incorporated herein by reference.

As indicated above, a common method of removing the iron present in theleachate is through the formation of a "jarosite" precipitate. Jarosite,MFe₃ (SO₄)₂ (OH)₆ where M is a monovalent ion, usually an alkali metal(general sodium or potassium) or ammonium, is commonly formed by addinga source of ammonium or sodium ions to the leach solution andmaintaining the solution at an appropriate pH by the addition of base.This process is also shown in FIG. 1. Various modifications to theso-called "jarosite process" are discussed in the article entitled "TheJarosite Process--Past, Present and Future", V. Arregui et al.,Lead-Zinc-Tin, TMS-AIME World Symposium on Metallurgy and EnvironmentalControl, 1980, J. M. Cigon, T. S. Mackey and T. J. O'Keefe, Ed., pp.97-123, incorporated herein by reference.

There are a number of problems associated with the formation of jarositewaste material. The jarosite can contain valuable metals such as silver,zinc, copper, lead, indium, etc. which require numerous expensiveprocess steps to recover. The jarosite can also contain toxic specieswhich can be leached into the environment by rain and groundwater.Therefore to avoid environmental contamination it is usually necessaryto store the jarosite wastes in sealed lagoons which are expensive tobuild.

A number of methods for treating such wastes have been disclosed.Steintveit et al. in Norwegian Patent 142,406 (1980) disclose a processfor leaching iron-containing waste with chloride-containing, acidicsolution at a temperature between 50° C. and the boiling point of thesolution. An alkali metal chloride or an alkaline earth metal chlorideis used as the source of the chloride with calcium chloride beingdisclosed as the preferred material. The pH of the solution is adjustedduring the leaching step so that the iron remains as a precipitate whilethe valuable metals are leached into the hot solution. The pH isadjusted to between about 2 and 4 preferably with calcium hydroxide. Itis disclosed that these conditions allow extraction of up to 95 percentof the lead and silver content of the waste. However, Applicants havefound that unpredictably with some jarosite wastes this process providesrecoveries of less than 20 percent of the silver present.

U.S. Pat. No. 4,054,638 of Dreulle et al. (1977) is directed to aprocess for recovering metals from sulfated residues from electrolyticzinc plants. The residue is digested preferably at a temperature between95° and 115° C. with hydrochloric acid in the presence of calciumchloride. This leaching process dissolves the metals present, includingthe iron, by forming the corresponding metal chlorides. Consequently,the process requires that the iron chloride be removed by extraction byan organic solvent. This process has a disadvantage of solubilizing theiron and requiring a separate separation step. There is no suggestion ordisclosure of using superatmospheric pressure for this leach.

U.S. Pat. No. 4,070,437 of Van Ceulen (1978), discloses a process forrecovery of metals from jarosite sludges. The process involves leachingthe jarosite with an acidic calcium chloride solution, preferably formedby mixing hydrochloric acid and calcium hydroxide or calcium carbonate.The leaching is preferably carried out close to the boiling point of theleaching medium. Insoluble calcium sulfate is formed and is separated byfiltration. This process has the disadvantage of solubilizingessentially all of the iron in the jarosite.

Another waste which contains metal values is zinc ferrite-containingmaterials. Modern electrolytic zinc processes commonly use a two-stepleaching process as depicted in FIG. 1. The second leaching stepinvolves a hot acid leach to dissolve zinc ferrite present. However,prior to the development of the two-step leach, a single neutral leachwas used which caused much of the zinc ferrite and associated metalvalues to be discarded as wastes. Therefore, there are existing wastelagoons which contain substantial quantities of zinc ferrite and othermetal values. The term "ferrite" is used herein to refer to a combinedmetal oxide-ferric oxide material, e.g. zinc ferrite (ZnO.Fe₂ O₃).

A number of processes have been developed for the purpose of recoveringthis zinc. One such process is disclosed by Rastas et al. in U.S. Pat.No. 3,959,437 (1976). Rastas et al. disclose a process in which theferrite of a non-ferrous metal, as well as the oxide of the non-ferrousmetal, is subjected to a neutral leach which dissolves most of the oxidebut leaves the ferrite substantially unaffected. The non-ferrous valuesin the solution are recovered and the undissolved ferrite material isfurther treated in a "conversion" stage with sulfuric acid-bearingsolution at atmospheric pressure and at a temperature of about 80° C. toabout 105° C. in the presence of alkali or ammonium ions. Under theseconditions, the non-ferrous metals dissolve as sulfates, while iron issimultaneously precipitated as an insoluble complex sulfate, i.e.,jarosite. U.S. Pat. No. 4,355,005 of Rastas et al. (1982), U.S. Pat. No.4,366,127 of Rastas et al. (1982), as well as U.S. Pat. No. 4,383,979 ofRastas et al. (1983) each disclose modifications to the processdisclosed in the '437 patent.

Steintveit in U.S. Pat. No. 3,684,490 (1972) discloses a method fortreating jarosite residue in which the residue is subjected to leachingwith sulfuric acid at a temperature of 50° to 95° C. and an acidconcentration of 10 to 70 grams per liter (hereinafter g/l). Theseleaching conditions are intended to decompose any zinc ferrites presentand provide for a greater recovery of the zinc.

U.S. Pat. No. 3,691,038 of Von Roepenack et al. (1972) discloses amethod for recovering zinc from oxides containing zinc and iron. Theoxide is leached with sulfuric acid at a temperature of 95° to 100° C.with an excess of sulfuric acid to solubilize the zinc and iron. Alkalimetal or ammonium ions are added to the liquid phase along with azinc-containing oxidic material at a temperature of 95° to 100° C. toprecipitate jarosite.

U.S. Pat. No. 4,192,852 of Pammenter et al. (1980) discloses a processfor treating zinc plant residues containing zinc ferrite andprecipitating the iron as a jarosite. The sulfate solution containingferric iron, free acid and non-ferrous metals is cooled, partiallyneutralized and then heated to a temperature not exceeding the boilingpoint at atmospheric pressure in the presence of sodium, potassium orammonium ions. U.S. Pat. No. 4,305,914 of Pammenter et al. (1981)discloses a process similar to that in the '852 patent.

U.S. Pat. No. 4,128,617, of DeGuire et al. (1978), describes athree-step process for the treatment of zinc calcine containing zincoxide, zinc sulfates, and zinc ferrites. The first step involves theneutral leaching of the zinc calcine with an effective amount of aqueoussulfuric acid containing solution. The leach residue is subjected to hotacid leaching with sulfuric acid followed by jarosite precipitation byalkali with the subsequent recycling of the jarosite-containing pulp.The preferred temperature range for the hot acid leaching is from about80° C. to the boiling point and preferably the temperature is greaterthan 90° C. There is no suggestion of the use of pressure.

The processes described hereinabove have one or more of thedisadvantages of (1) having low rates of extraction of metal values, (2)providing low levels of recovery of certain metal values, (3) havingpoor filter-ability of the iron-containing residue, and/or (4)solubilizing large amounts of iron.

Several processes disclosed in the art have used elevated temperatureand pressure leaching steps in treating zinc plant residues.

U.S. Pat. No. 3,143,486 of Pickering et al. (1964) discloses a processfor the extraction of zinc from zinc ferrite containing residue. Theprocess involves subjecting the residue to a first-stage leachingtreatment under non-oxidizing conditions in a closed vessel in thepresence of excess sulfuric acid at a temperature between 140° C. and260° C. Zinc is dissolved as well as ferrous sulfate which is stable atthe temperatures and acidities used. Ferric iron is precipitated as abasic sulfate. The leachate is then subjected to a second-stage leachingtreatment at 140° C. to 260° C. under oxidizing conditions to oxidizethe ferrous sulfate to ferric and precipitate the ferric material.Similarly, U.S. Pat. No. 3,493,365 of Pickering et al. (1970) disclosesa two-step high temperature method of treating zinc plant residuecontaining zinc ferrite. This process differs from that of the '486patent in that in the second step a source of a cation selected from thegroup consisting of sodium, potassium and ammonium is added in order toprecipitate the iron from the liquor as a jarosite material.

A process for treating sulfide ores which involves a two-step leach isdisclosed by U.S. Pat. No. 4,266,972 of Redodno-Abad et al., (1981). Thefirst leach uses sulfuric acid under an oxygen atmosphere at 150° to250° C. Zinc and copper are solubilized with lead, the noble metals, andiron oxide remaining as a residue. After a solid liquid separation, thefiltrate is adjusted to a pH of 1.5 to 2. Sodium chloride, calciumchloride and ferric chloride are added to precipitate calcium sulfate.The leach is conducted at a temperature between 60° C. and 90° C. withthe silver, lead, and gold being solubilized as the chlorides. The ironoxide remains as a residue. After a solid/liquid separation, the silver,lead, and gold are recovered by cementation with zinc, with the liquidbeing subjected to an extraction to recover the zinc.

None of these processes which use high temperature and pressurizedleaches discloses or suggests that jarosite-containing wastes can beadvantageously treated in such a manner. In fact, most disclose the useof an oxidizing atmosphere to form ferric iron which will precipitate.These patents also disclose that potassium ions can be added to a zincferrite leach solution in order to precipitate potassium jarosite. Asdiscussed in detail hereinbelow, it has been found that the recovery ofmetal values can be unpredictably affected by the presence of potassiumions.

Accordingly, there is a need for a process to treat jarosite and ferritecontaining wastes from zinc recovery processes in order to recover metalvalues which are contained in the waste materials and render the residuesuitable for disposal as a nonhazardous waste. There is also the needfor a process which will not be subject to the unpredictable effect ofpotassium ions.

SUMMARY OF THE INVENTION

It has now been found that the above described disadvantages of knownprocesses can be minimized or eliminated by the instant invention.According to the present invention, a process is provided for recoveringmetal values from jarosite wastes from a zinc recovery plant saidprocess comprising leaching the waste with an acidic solution of calciumchloride at or above the atmospheric boiling point of the solution andunder at least the atmospheric pressure.

One of the embodiments of the instant invention comprises a process forrecovering metal values from jarosite-containing wastes from a zincplant using an acidic solution of calcium chloride at a temperatureabove the boiling point of the solution and under super-atmosphericpressure.

Another embodiment of the instant invention comprises a process forrecovering metal values from a jarosite-containing waste wherein theprocess comprises contacting the waste at a temperature of between about110° C. and 300° C. and a pressure of at least the auto genous pressurewith a solution containing between 1.0 and 5.0 molar calcium chloride.The contacting solution has a pH of between about 2.0 and 3.5. This pHis maintained by the addition of a calcium compound selected from thegroup consisting of calcium oxide, calcium hydroxide, calcium carbonateand mixtures thereof.

In a further embodiment, the instant invention comprises a process inwhich jarosite-containing wastes are leached with an acidic solution ofcalcium chloride at greater than atmospheric pressure and a temperaturegreater than the atmospheric boiling point of the solution to form aliquid leachate and a solid residue. The liquid leachate is separatedfrom the solid residue and contacted with a reducing metal to reducesilver cations contained in the leachate to metallic silver. Themetallic silver is then separated from the liquid solution. The liquidsolution is sulfided by mixing the solution with a sulfide compound toprecipitate the lead contained in the solution as lead sulfide. Thesolid lead sulfide is separated from the liquid phase. Substantially allof the zinc in the remaining liquid phase is recovered using a zincrecovery process to provide a liquid solution substantially free ofzinc. The resulting liquid solution is heated with calcium oxide toprovide a vapor containing ammonium hydroxide and bottoms which containcalcium chloride.

In another embodiment, the instant invention comprises contacting aferrite which contains metal values with a leach solution of sulfuricacid and ammonium ions. The leach solution containing the ferrite isheated at a temperature above about 90° C. for up to about 12 hours toform a solid containing ammonium jarosite and a liquid phase. The solidphase is mixed with an acidic leaching solution containing calciumchloride. This mixture is heated at superatmospheric pressure and at atemperature above the atmospheric boiling point of the solution tosolubilize a substantial portion of the metal values.

BRIEF DESCRIPTION OF THE DRAWING

FIG. 1 shows a typical process flowsheet for an electrolytic zinc plant;

FIG. 2 shows the effect of potassium on metals extraction;

FIG. 3 shows the effect of temperature and potassium concentration onsilver extraction;

FIG. 4 shows the flowsheet for a high temperature calcium chlorideleach;

FIG. 5 shows a process flowsheet for the recovery of metal values fromjarosite;

FIG. 6 shows a process flowsheet for a ferrite leach followed by a hightemperature calcium chloride leach;

FIG. 7 shows a process flowsheet for metals recovery from ferrite andjarosite in association with a zinc plant; and

FIG. 8 shows a process flowsheet of a preferred embodiment.

DESCRIPTION OF THE PREFERRED EMBODIMENTS

The process of this invention comprises a method for recovering metalvalues from jarosite containing wastes from electrolytic metal recoveryplants. The jarosite waste is leached at or above the boiling point ofthe leach solution under pressure with an acidic calcium chloridesolution. Optionally, non-jarosite wastes such as zinc ferrite can besubjected to a preleach step in which the iron is converted to jarositematerial and the zinc is substantially solubilized.

The use of a high temperature, pressurized leaching process for leachingjarosite has been found to provide several unexpected advantages overthe lower temperature, atmospheric pressure processes described in theprior art. It has been found that the final leach slurry unexpectedlyfilters two to three times faster than comparable slurries obtained witha 95° C. to 100° C. leach such as that described by Steintveit inNorwegian Patent No. 142,406 (Supra). Additionally, the leach times aresignificantly less than those required when leach temperatures below100° C. are used. This procedure has also been found to provide higherextractions of metal values from certain types of wastes than obtainedwith the lower temperature leaches.

It has also been found that certain materials unexpectedly interferewith the recovery of metal values from jarosite. In particular, it hasbeen observed that with some jarosite wastes the presence of potassiumions can result in significantly lower recoveries of metal values suchas silver and lead from the leach solution. FIG. 2 is a graphicpresentation of experimental data showing the effect of potassium on theamount of metal extracted. This effect is particularly evident at lowerleaching temperatures. Thus, using higher leaching temperatures thandisclosed in the prior art appears to unexpectedly compensate for thepresence of potassium in the leaching solution. Graphic evidence of thispreviously unrecognized phenomenon is provided in FIG. 3 where theeffect of increasing the extraction temperature from 130° C. to 180° C.is shown. The results also show that only 10 percent of the silver wasextracted at 95° C. when 5 grams per liter of potassium were present.With 10 grams per liter of potassium, 80 percent of the silver wasextracted at 180° C.

The detrimental effect of ionic potassium is not observed with everyjarosite material. Differences in the effect of potassium concentrationon the recovery of silver have even been observed on samples of jarositeobtained from different locations in the same waste lagoon. Althoughextensive work has been done to determine the basis of this phenomenon,the reason remains unknown. A method for predicting the magnitude ofthis effect has not been discovered as yet. Therefore, practicing thepresent invention assures that the recovery of metal values can bemaximized even with variations in the magnitude of the potassium effect.

The leaching process for ammonium jarosite can be represented by thefollowing reaction scheme.

    2NH.sub.4 Fe.sub.3 (SO.sub.4).sub.2 (OH).sub.6 +CaCl.sub.2 +3Ca(OH).sub.2 →6Fe(OH).sub.3 +2NH.sub.4 Cl+4CaSO.sub.4           (1)

The overall leach process, including regeneration of CaCl₂, may berepresented by combining the above reaction with the reaction occurringduring the "lime boil" operation,

    2NH.sub.4 Cl+Ca(OH).sub.2 →CaCl.sub.2 +2NH.sub.3 +2H.sub.2 O (2)

to get the following overall reaction,

    2NH.sub.4 Fe.sub.3 (SO.sub.4).sub.2 (OH).sub.6 +4Ca(OH).sub.2 →6Fe(OH).sub.3 +4CaSO.sub.4 +2NH.sub.3 +2H.sub.2 O. (3)

The calcium chloride leaching process of the instant invention isnormally conducted at or in excess of the atmospheric boiling point ofthe leach solution. As used herein, the terms "atmospheric boilingpoint" and "boiling point" are used interchangeably to refer to thetemperature at which the solution boils under the particular atmosphericpressure to which it is being subjected. Thus the boiling point for asolution of the same composition can vary depending upon thegeographical elevation at which it is heated. Ordinarily, the boilingpoint of the leach solution is in excess of 110° C. It has been foundthat leaching at a temperature of at least the boiling point of thesolution and preferably above 110° C. provides improved results inextracting metal values from the jarosite-containing waste materials.The boiling point of the leaching solution can be increased byincreasing the concentration of calcium chloride in the solution.Therefore, it is contemplated that the instant invention encompasses theembodiment of heating a solution at or above 110° C. even though theboiling point of the solution is in excess of 110° C. Leachingtemperatures from the boiling point to about 300° C. can be used withtemperatures of between about 120° C. and 225° C. being preferred andtemperatures of between about 130° C. and 200° C. being most preferred.The pressure on the leaching solution has not been found to be criticalas long as the pressure is sufficient to prevent, the solution fromboiling. Ordinarily, autogenous pressure is maintained in the system,i.e. greater than 1 atmosphere and preferably at least about 1.1atmospheres. As used herein, the term "autogenous pressure" means thatpressure which develops in a closed system when the system is heated.

The concentration of calcium chloride should be maintained above about0.5 molar in the leaching solution to achieve the best extraction andsolubilization of metal values. The concentration can range from about0.5 molar up to the saturation point for calcium chloride for theparticular temperature and solution. It is preferred that theconcentration of calcium chloride be between about 1 molar and about 4molar in the leaching solution with the most preferred concentration ofcalcium chloride between about 2 and 3.5 molar.

Calcium chloride can be added to the leaching solution directly or canbe formed in situ by using a chloride source and a calcium source. Forexample, materials such as hydrogen chloride, sodium chloride andammonium chloride can be used as the chloride source. Useful sources ofcalcium include lime, hydrated lime, and calcium carbonate. Theparticular materials used depends upon the economics. The use of thecalcium oxide or calcium hydroxide as the calcium source has theadditional advantage that the pH of the solution can be adjusted to thedesired range by their addition. To avoid the buildup of certaincations, e.g. sodium, in a continuous system, it may be necessary tobleed a small stream out of the recycle or to remove the cation by othermeans such as ion exchange.

For effective leaching with calcium chloride, it is important that thepH of the solution be less than about 5. It is preferred that the pH beless than about 4 and most preferred that the pH be less than about 3.For best results, the extraction should be carried in a pH range ofabout 2.0 to 3.0 During the calcium chloride leach of the jarosite, thepH of the solution can decrease to a pH of less than 1.0 unless adjustedby the addition of a base. Although any base, such as sodium hydroxideor sodium carbonate can be used, it is advantageous as discussedhereinabove to use a calcium containing base such as calcium oxide,calcium hydroxide or calcium carbonate.

Other things being equal, the higher the leaching temperature, theshorter the leaching time required to achieve a particular level ofextraction of metal values. The time required for the extraction dependsupon the concentration of calcium chloride in the leach solution, thetemperature at which the leaching is conducted, the pH of the solutionand the particular waste feed being treated. Ordinarily, the jarositeleaching requires less than about 6 hours with leach times in the rangeof about 5 minutes up to about 2 hours being preferred. At the higherleaching temperatures, it is expected that in excess of about 75 percentand preferably in excess of 85 percent of the silver and lead present inthe jarosite is solubilized during the preferred leaching time.

The jarosite-containing waste materials suitable for use as feed to thecalcium chloride leaching process comprise materials which contain themetal values such as zinc, silver, lead, indium. The preferred feeds arejarosite residues from an electrolytic zinc process. The material calledjarosite has the formula MFe₃ (SO₄)₂ (OH)₆. There are a variety ofjarosite type compounds which contain different ions in place of the M,for example, potassium, lead, silver, sodium, rubidium or ammonium.Ammonium jarosite is a principal effluent from an electrolytic zincprocess.

Other iron containing materials can be used as feeds to produce jarositee.g. zinc ferrite (ZnFe₂ O₄). The zinc ferrite used can be any suchmaterial containing metal values. The source of zinc ferrite isordinarily waste lagoons from zinc plants.

An alternative source of zinc ferrite material is the dust formed duringsteelmaking in an electric arc furnace (EAF). The dust primarilycomprises zinc ferrite, zinc oxide, and various forms of ferric oxide.The dust can also contain lead, cadmium and chromium and is thereforeusually classified as a hazardous waste. The dust can additionallycontain metal values such as silver. Thus, the dust can be fed into theprocess of the instant invention to allow recovery of metal valuesand/or removal of the toxic metals.

In order to be used in the present process, the EAF ferrite is subjectedto a leach operation in which the ferrite is contacted with sulfuricacid and an ammonium source to precipitate ammonium jarcsite whichcontains some of the metal values. The ammonium jarosite can then beconveyed to the calcium chloride leaching process.

In the practice of the instant leaching process, the jarosite 5 is addedto the leaching vessel 6 along with calcium chloride as represented inFIG. 4. The calcium chloride can be makeup or be recycled fromdownstream process steps or a mixture of the two sources. During thehigh temperature leaching process, the pH of the solution tends todecrease due to generation of acid. The pH is adjusted to the preferredleaching range by adding a basic material, preferably calcium oxide,calcium carbonate, and/or calcium hydroxide. The solution is heated tothe leaching temperature, preferably by steam, with agitation of themixture. After the appropriate leaching time, the resulting slurry issubjected to a liquid-solid separation 7 to provide solid tails 8 and aliquid leachate 9. The solid tails contain iron oxides, gypsum, and/orunreacted gangue while the liquid leachate contains solubilized zinc andother metal values. The tails are discarded while the leachate issubjected to further downstream treatment as discussed in more detailhereinbelow. Optionally, a portion of the leachate is recycled 10 to thecalcium chloride leach step to effect a build-up of metals in thepregnant solution.

The leachate from the calcium chloride leaching process can be treatedby any suitable method to remove the metal values such as copper,silver, or gold which might be present. When silver and copper arepresent, it is preferred that a metal cementation process be used toremove the silver and/or copper. As represented in FIG. 5, this can beaccomplished by treating the leachate with any material which is asuitable reductant for the metal values. Metals which have higherreduction potentials than silver, such as lead or zinc, can be used toform a silver cement. Presently for economic reasons, it is preferred touse zinc dust. Zinc dust can be added to precipitate a solid consistingof a mixture of silver, lead, copper (depending upon the presence of theparticular metal in the solution) and zinc. The solid product which isseparated from the liquid phase can then be subjected to additionalstandard metallurgical processes to recover the pure metal values. Theseparation of the solid product from the liquid phase can beaccomplished by standard solid-liquid separation techniques such asfiltration or centrifugation.

The liquid from the metal cementation process contains dissolvedreducing metal, calcium chloride, lead, zinc and other noncementableimpurities. The liquid can then be subjected to additional processingsteps to separate the reducing metal and other metals such as lead andzinc (if not used as the reducing metal) from the liquid. Optionally,the liquid can be subjected to a sulfide precipitation procedure torecover the lead and any trace of heavy metals which precipitate assulfides under these conditions. The sulfide precipitation can beaccomplished by adding a sulfide-containing material such as hydrogensulfide to the liquid to form solid lead sulfide. The liquid can then beseparated from the solid lead sulfide cake by conventional solid-liquidseparation techniques. Alternatively, the pH of the liquid can beadjusted to precipitate metals such as lead and zinc as theirhydroxides. For example, calcium oxide can be added to precipitate leadhydroxide and zinc hydroxide. In another embodiment, the liquid from themetal cementation process is subjected to an extraction process toseparate zinc as described herein below. The raffinate can be partiallyneutralized to a pH of about 8-9 to form a lead hydroxide precipitatewhich is separated from the liquid. The liquid can then be subjected toa lime boil as discussed below. The particular choice of procedures tobe used to treat the liquid will depend upon the metals present, theirconcentrations and the economics of the various alternatives. The solidsrecovered can be disposed of in an appropriate manner or can besubjected to additional processing steps to recover any metal valuespresent.

The liquid from the lead removal process can be subjected to zincrecovery procedures. Any standard method of recovery suitable forremoving zinc from such a solution can be used. A preferred method issolvent extraction in which the aqueous solution containing zinc iscontacted with an extractant which transfers the zinc from the aqueoussolution to the organic phase. The second phase can be a solvent whichis not miscible with the aqueous solution or it can comprise theextractant. Alternatively, the liquor containing the zinc can be passedthrough an ion exchange resin to remove the zinc. A preferred methodinvolves the use of extractants such as di-2-ethylhexyl phosphoric aciddissolved in a hydrocarbon solvent such as kerosene. In the use of suchmaterials, the pH is adjusted to determine the phase in which the zincultimately resides. In a preferred method of operation, the zinc speciesis selectively extracted into the organic phase by maintaining the pH inan appropriate range depending on the extractant used. The pH can beadjusted with a base such as calcium oxide, calcium hydroxide, calciumcarbonate, etc. Zinc is then stripped from the loaded organic with astrong acid solution such as spent electrolyte from a zinc plant. Stripliquor containing the zinc can be transferred to the zinc plant forelectrowinning.

The raffinate from which the zinc has been removed can be subjected toadditional processing to recover the calcium chloride and ammonia. Abase can be added to increase the pH of the solution to a basic range,preferably above about 9. Although bases such as NaOH or KOH can beused, it is preferred that a basic calcium compound such as calciumoxide, calcium hydroxide, calcium carbonate or mixtures thereof be usedto avoid build-up of undesired materials such as potassium ions in thesystem if recycle is used. Preferably a "lime boil" operation can beused in which lime is added to the raffinate to increase the pH to abasic range, preferably above about 9, and the mixture heated to formammonia which can be removed by volatilization. Preferably, atemperature of at least the boiling point of the solution is used. Theammonia formed is removed, usually in the vapor, and can be used to formjarosite in the ferrite treating process described hereinbelow or can betransferred to the zinc plant. The brine solution from the lime boilprocess contains calcium chloride as well as residual metal values. Thisbrine solution can be recycled for use in the calcium chloride leachingprocess.

An alternative method for recovering the zinc is to treat the liquorfrom the sulfiding procedure with a base such as calcium oxide toprecipitate zinc hydroxide. In this operation (not shown in FIG. 5),sufficient base is added to provide a solution pH of between about 8 and10. Bases which are useful in this process include calcium hydroxide,calcium oxide, sodium hydroxide, etc., provided there is no detrimentalbuild-up of cations in the system. The zinc hydroxide precipitate isseparated from the liquor by standard liquid-solid separation techniquessuch as filtration or centrifugation. The zinc hydroxide solid ispreferably transferred to a zinc plant for recovery of the zinc. Theliquor from this precipitation process is transferred to a lime boilprocess as described hereinabove.

Depending upon the level of lead and/or zinc present in the liquor fromthe silver cementation process, the basic precipitation step can beoptionally performed on the liquor from the silver cementation step.This will result in a mixture of lead hydroxide and zinc hydroxidesolids being formed. After a solid/liquid separation, these solids canbe disposed of in any manner appropriate. The liquor from thisprecipitation step is transferred to a lime boil process as describedhereinabove.

When zinc ferrite (ZnFe₂ O₄) is a significant component of the waste, asulfuric acid (sulfuric) leach step is used. In this leach there is asimultaneous leach of the ferrite and a precipitation of the iron in theform of jarosite. A method for this simultaneous leach and precipitationis described by Rastas et al. in U.S. Pat. No. 3,959,437 (1976). Rastas,however, is only concerned with solubilizing the zinc in the ferrite anddoes not treat the jarosite to recover metal values.

In the present invention, the ferrite material is combined with sulfuricacid and an ammonium source at a temperature above about 50° C. andbelow the decomposition temperature of ammonium jarosite. Preferably,the temperature is between about 80° C. and about 170° C. The sulfuricacid used is preferably spent electrolyte from an electrolytic zincplant and is in the ferrite leach solution to the extent of betweenabout 10 and about 60 grams per liter, preferably between about 30 andabout 50 g/l. The ammonium compound is preferably recovered from thelime boil process step described hereinabove. This leach is ordinarilyconducted at a temperature of about 80° C. to 100° C. However, asdiscussed hereinabove, the leach can be accomplished at a temperatureabove the boiling point of the solution under pressure. It iscontemplated that the leach is conducted under autogenous pressure atthe elevated leaching temperatures.

The ferrite leach is conducted for a period sufficient to precipitatejarosite and solubilize zinc. Preferably the leach time is about 12 to36 hours. Commonly, at least about 70 percent of the zinc present in theferrite is solublized with less than five percent of the iron andessentially no silver or lead solubilized. The solid jarosite containingmetal values and any remaining untreated ferrite is separated from theliquid phase by standard solid-liquid separation techniques such asthickening, filtration or centrifugation. The liquor from the ferriteleach step which contains zinc sulfate is preferably conveyed to a zincplant for recovery of the zinc. The solid jarosite is subjected to thecalcium chloride leach process described hereinabove.

Referring now to FIG. 4, a process involving a calcium chloride leach ofjarosite in a system at a temperature above the boiling point of thesolution is represented. Preferred embodiments cf this process arehereinafter described. Sufficient jarosite 5 is added to the leachingvessel 6 to provide a slurry containing about 10 to about 40 weightpercent jarosite. A chloride source such as calcium chloride is added tothe slurry. Alternatively, other chloride sources can be used dependingupon the economics of the process. The pH is maintained in the desiredrange of about 1.5 to about 3.5 by the addition of lime, i.e. calciumoxide, and/or calcium hydroxide, calcium carbonate. The slurry is heatedto a temperature ordinarily above the boiling point of the solution,normally above 110° C. and preferably above about 120° C. The leachingzone or leaching vessel must be capable of withstanding the equilibriumpressure at the temperature selected. While the vessel can bepressurized with an inert gas, in ordinary operation it is maintainedunder autogenous pressure, i.e. the pressure established in the vesselby the vaporization of volatile components such as water at the leachingtemperature. Ordinarily, the pressure in the leaching zone will bebetween about 5 to about 210 psig with the preferred range being fromabout 15 to about 130 psig.

The leaching operation can be carried out in either a batch or acontinuous mode. The particular choice of operation will depend upon theleaching time necessary to extract the desired amount of the metalvalues. In a batch operation, the necessary quantity of calcium oxide,calcium carbonate and/or calcium hydroxide can be introduced initiallyinstead of being added incrementally during the leaching process. Theappropriate amount of base can be readily calculated by one skilled inthe art based upon the jarosite content of the feed material added tothe leach. The initial quantitative addition of the calcium oxide isoperationally advantageous since it eliminates the equipment required toadd the material incrementally throughout the leaching process as wellas the need to monitor the pH of the system. We have found that there isno difference in the final level of the metal values extracted when thecalcium oxide is added initially as opposed to incrementally during theleaching operation as long as the final pH of the leaching solution iswithin the desired range.

At the conclusion of the leaching process, the liquid-solid slurry isseparated 7 into solid tails 8 and a liquid leachate 9. This can beaccomplished by any known method for solid-liquid separations, althoughwith this type of slurry, it is preferred that the separation beaccomplished by filtration accompanied by appropriate washings.Surprisingly, it has been found that slurries from the high temperatureleaches filter two to three times faster than the slurries from a leachconducted at 95° C. to 100° C. This is an unexpected advantage ofoperating at the higher leach temperatures. The solid tails whichcontain gypsum and iron oxides are discarded. Optionally, a slip streamfrom the leachate is recycled 10 to the calcium chloride leachingprocess to allow further concentration of metal values. The remainingleachate liquid is subjected to further processing for recovery of metalvalues.

The recovery of the metal values can be accomplished by any known methoduseful for separating such materials. A particular separation scheme isset forth in FIG. 5. The leachate 9 from the calcium chloride leach 6 iscontacted 13 with a metal 14 capable of reducing silver. Preferably zincmetal is used although lead is also suitable and the choice is generallya matter of economics. The metal, ordinarily in the form of fineparticulates, is contacted with the leachate preferably at a temperaturebetween about 40° C. and about 80° C. The silver cement 15 is separated17 from the liquid by filtration although centrifugation could also beused. The silver is recovered from the silver cement by conventionalmeans such as smelting.

As shown in FIG. 7, lead can be separated from the liquor from thecementation process by sulfiding 51 with hydrogen sulfide or anyappropriate sulfide source to precipitate lead sulfide 53. Ordinarily,greater than about 95 weight percent of the lead is precipitated. Thelead sulfide is separated from the liquid by filtration althoughcentrifugation can also be used.

The liquid from the sulfiding step is treated 19 to recover zinc.Solvent extraction and precipitation are preferred methods of separationalthough any method suitable for recovering zinc 21 from such a streamcan be used. The zinc-solvent extraction 55 process, depicted in FIG. 7and FIG. 8, involves contacting the aqueous solution containing the zincwith an extractant which will bind the zinc and allow its extractioninto an organic phase. A preferred extractant is di-2-ethylhexylphosphoric acid dissolved in an appropriate diluent. In operation, theextractant is mixed with the zinc feed solution. Since the extraction ofzinc lowers the aqueous pH, a base such as NaOH, CaO, or CaCO₃ can beadded to maintain the desired pH. At a pH of between about 2 and 3 thezinc is selectively extracted into an organic phase 56 in contact withthe aqueous solution. This organic phase containing the zinc isseparated from the aqueous phase and the zinc is then stripped 57 into astrong sulfuric acid solution. The solution 59 containing the zinc canbe transported to a zinc plant for recovery of the zinc.

In another process sequence (shown as a part of the scheme in FIG. 8),the liquid from the cementation step is subjected to solvent extractionto remove the zinc as described hereinabove. The raffinate liquor fromthe zinc solvent extraction is partially neutralized, preferably withcalcium oxide, to a pH of between about 8 and 9 to precipitate leadhydroxide. The lead hydroxide solid is removed, for example byfiltration 65 or centrifugation and disposed of or the lead can berecovered by known procedures.

The liquid 23 recovered from this separation is subjected to a lime boil25 by adding lime to provide a pH above about 9 and heating to about theboiling point of the mixture. Ammonia is vaporized and passed through acondensor to provide an ammonium hydroxide stream which can be recycledto the ferrite leach or to the zinc plant to produce ammonium jarosite.The liquid condensate can be neutralized with sulfuric acid to produceammonium sulfate suitable for fertilizer.

Alternatively (not shown), the zinc can be precipitated as zinchydroxide by adding calcium oxide to the liquor from the sulfidingprocess. Sufficient calcium oxide is added to provide a solution pH inthe range of about 8 to 10. The precipitated zinc hydroxide is separatedby liquid-solid separation means preferably filtration. The recoveredzinc hydroxide can be conveyed to a zinc plant for recovery of the zinc.

As shown in FIG. 8, the liquor from the zinc separation process issubjected to a lime boil treatment with calcium oxide. This lime boilstep is conducted at a temperature effective to form ammonia and calciumchloride commonly above room temperature. Sufficient calcium oxideand/or calcium hydroxide is added to provide a pH of above about 9. Thisprocess step provides an ammonia overhead vapor stream and a brinestream containing calcium chloride and/or a precipitate of leadhydroxide. The brine stream is preferably recycled to the calciumchloride leach step to allow reuse of the chloride values. The ammonia,which can be dissolved in water or used directly, is preferably recycledto a zinc plant or to a ferrite preleach process.

In an alternative procedure (not shown), the liquor from the cementationstep can be combined with sufficient calcium oxide to cause the combinedprecipitation of zinc hydroxide and lead hydroxide. Calcium oxide isadded to provide a pH in the range of about 8 to 10. The metalhydroxides are separated by usual solid-liquid separation techniquespreferably by filtration. These hydroxides can then be returned to azinc plant for recovery of the zinc. The liquor from the precipitationprocess is then subjected to a lime boil as discussed hereinabove. Theresulting brine solution is preferably recycled to the calcium chlorideleach with the ammonium hydroxide formed preferably being recycled to azinc plant or sold as a by-product.

When zinc ferrite is treated, a sulfuric acid ferrite leach operation isused as discussed hereinabove. As shown in FIG. 6, an aqueous slurry ofthe ferrite solids 31 is combined with sulfuric acid and ammoniumhydroxide. Sufficient ferrite solids are added to provide a slurrycontaining about 30 weight percent ferrite. Sulfuric acid is added tothe extent necessary to provide a final acid level of about 10 to about60 grams per liter. Preferably, the sulfuric acid used is spentelectrolyte from a zinc plant. Sufficient ammonia is added to accomplishprecipitation of the ammonium jarosite. Suitable ammonium sourcesinclude without limitation ammonia, ammonia water, ammonium hydroxideand ammonium salts such as ammonium sulfate. Preferably, the ammonia isreceived from the lime boil operation described hereinabove. Preferably,ammonium ions are present at the end of the leach to the extent of about10 g/l to about 20 g/l of the solution. About 15 grams of ammonia perliter of slurry is used in this operation. The ferrite leach 33 isaccomplished at a temperature of about 95° C. for a period of about 12to 24 hours. This resulting slurry is subjected to a thickeningliquid-solid separation procedure. The addition of a small amount of"seed" jarosite can be added to aid in the formation of solid jarosite.The liquid 35 from this ferrite leach is rich in zinc and can beconveyed to a zinc recovery process. The solid material 37 ispredominantly jarosite which contains zinc and other metal values. Thismaterial is conveyed to a calcium chloride leach 6 as describedhereinabove.

In FIG. 7 there is shown an embodiment of the instant invention in whichthe product 41 from a ferrite leach 33 is combined with additionaljarosite 43 in a repulp tank 45 and the resulting pulp 47 is fed to acalcium chloride leach 6. The subsequent metal recovery steps have beendescribed in detail hereinabove. Also shown in this embodiment is themovement of various streams to and from a zinc plant 49. The zinc-richliquid 35 from the ferrite leach is transferred to the zinc plant forrecovery of the zinc. Also, the strip liquor from the zinc extraction istransferred to the zinc plant. Ammonia from the lime boil is used in thezinc plant and/or the ferrite leach step to form jarosite. The spentsulfuric acid electrolyte from the zinc plant is used in the ferriteleach and in zinc stripping. As discussed hereinabove, an alternativemethod of zinc separation which is not shown is the precipitation ofzinc hydroxide. This zinc hydroxide can also be conveyed to the zincplant for recovery of the zinc metal.

In FIG. 8 is shown a preferred process scheme for the treatment offerrite and/or jarosite wastes. A ferrite containing feed 31 issubjected to a leach 33 by contacting it with sulfuric acid and anammonium source as described hereinabove. The source of the sulfuricacid is spent electrolyte 71 from a zinc plant 49. The ammonium sourceis ammonia or ammonium hydroxide 73 recycled from the lime boil 25. Aflocculant 75 is added to the slurry and the solids are separated fromthe liquid. The liquid 77, containing zinc, is conveyed to a zinc plantfor zinc recovery. The solids, containing ammonium jarosite and metalvalues, are transferred to a repulp tank 45 where these solids can bemixed with additional jarosite feed 43 and slurried with a calciumchloride source. The leach mixture 6 is heated to above 110° C. underpressure and stirred. The pH of the leach is adjusted to the desiredlevel by the addition of calcium oxide, calcium carbonate and/or calciumhydroxide. After leaching, the slurry is filtered 79 with washing toseparate the solid tails 8 and the leachate 9.

The leachate is contacted 13 with metallic zinc to form a silver/coppercement 15. The cement is removed, preferably by filtration 81, and thesilver and copper are purified by standard metallurigical methods. Theliquid is mixed 55 with an extractant with the pH of the solutionadjusted by adding calcium carbonate. The extractant binds the zinc andallows its extraction into an organic phase. The phases are separatedand the zinc-containing organic phase 56 is contacted 57 with spentelectrolyte 71 from a zinc plant to strip the zinc from the organicphase into an aqueous phase. The aqueous phase 59 containing the zinc istransferred to a zinc plant for recovery of the zinc. The organic phase83 containing the barren extractant is recycled to contact freshzinc-containing solution.

The aqueous phase from which substantially all of the zinc has beenremoved is contacted with lime 85 to increase the pH to between about 8and 9 to precipitate lead hydroxide. The solid lead hydroxide isseparated from the liquid by filtration 65. The liquid phase is mixedwith lime and heated to boiling. The liquid from the lime boil 25 is acalcium chloride-brine solution and is recycled 87 to the calciumchloride leaching step. The vapor containing ammonia is condensed 89with cold water and the resulting ammonium hydroxide solution isrecycled 73 to the zinc plant and/or to the ferrite leach. Anyuncondensed vapor can be passed into an acid scrubber using, forexample, a sulfuric acid wash to form a solution of ammonium salt suchas ammonium sulfate.

The following examples are given for illustrative purposes only and arenot to be a limitation on the subject invention.

EXAMPLE 1

Three jarosite feed materials were used in the Examples. Sample 1 wastaken from a first zinc plant waste lagoon. Samples 2 and 3 were takenfrom different locations in a second zinc plant waste lagoon. The assayresults are given in Table 1A.

                  TABLE 1A                                                        ______________________________________                                        Element, % Sample 1     Sample 2 Sample 3                                     ______________________________________                                        Zn         4.95         7.53     7.46                                         Ag oz/T    7.53         12.3     11.2                                         Fe         27.1         22.6     19.9                                         Pb         2.13         6.98     7.10                                         NH.sub.4   2.04         1.01     1.20                                         K          0.054        0.364    0.337                                        Na         0.122        0.162    0.156                                        Cu         0.32         0.3      0.347                                        Mn         0.362        5.17     4.47                                         In         0.010        --       0.008                                        ______________________________________                                    

A series of runs were made in which the jarosite feed material wascontacted with a leaching solution containing 330 grams per litercalcium chloride, for 5 hours unless otherwise noted. The pH wasmaintained in the range of 1.8-3.5 by the addition of calcium hydroxide.The amount of solids in the leach was about 16.8 weight percent.

The leaching process was conducted at different temperatures asindicated in Table 1B. When the temperature of the leach solution was ator below the boiling point of the solution, the leach was conducted inconventional glassware. When the leaching temperature was above theboiling point of the solution, the leach was conducted in a 2-liter Parragitated autoclave.

The leachate was separated from solids by filtration through a Buchnerfunnel using low vacuum.

                  TABLE 1B                                                        ______________________________________                                                  Leach                                                               Run       Temperature,                                                                             Extraction, %                                            No.       °C. Ag        Pb   Zn                                        ______________________________________                                        Sample 1 Jarosite (5-10 g/l K)                                                1         95         76        81   49                                        2         103 (boiling)                                                                            78        77   47                                        3         110        93        89   44                                        4         120        96        85   42                                        5         130        99        94   48                                        Sample 3 Jarosite (25 g/l K)                                                  6         95         72        84   28                                        7         110        72        (62) 28                                        8         120        73        (62) 29                                        9         130        73        88   29                                        10        140        72        89   31                                        11        150        74        80   30                                        12        180 (1 hr) 76        86   37                                        13        200 (2 hr) 75        84   34                                        14        225 (1 hr) 75        87   35                                        15.sup.(a)                                                                              325        74        80   31                                        ______________________________________                                         .sup.(a) Run in the nickel bomb reactor for approximately 5 minutes. Bomb     leach extractions are generally lower than comparable autoclave values.       Description of nickel bomb reactor in Example 3.                         

EXAMPLE 2

The effect of potassium on the recovery of metal values from certainfeeds is shown in Table 2. The leaching conditions and apparatus ofExample 1 were used. The conventional glassware was used for the 95° C.to 103° C. leaches with the Parr autoclave used for the leachesconducted at temperatures greater than 103° C. With Sample 2 feed, therecovery of metal values was significantly reduced when potassium waspresent. Increasing the leach temperature in Run 18 however, allowedrecovery of the metal values even in the presence of potassium. Sample 3feed showed no potassium effect.

                  TABLE 2                                                         ______________________________________                                        Conditions:                                                                              95-103° C. tests:                                                                   330 g/l CaCl.sub.2 ; 4-5 hours;                                               pH 1.8-3.5.                                                      >103° C. tests:                                                                     autoclave; 330 g/l CaCl.sub.2 ;                                               1 hour; pH 1.8-3.5.                                   ______________________________________                                                      Leach      K in Leach                                           Sample                                                                              Run     Temp       Solution                                                                              Extraction, %                                No.   No.     °C. g/l     Ag   Pb   Zn                                 ______________________________________                                        1     1       95         5-10    76   81   49                                       2       103 (boiling)                                                                            5-10    78   77   47                                       3       110        5-10    93   89   44                                       4       120        5-10    96   85   42                                       5       130        5-10    99   94   48                                 2     16      95-100     0       72   84   41                                       17      95-100     5       13   35   27                                       18      180        10      79   88   (60)                               3     19      95-100     0       73   86   34                                       6       95-100     25      72   84   28                                       12      180        25      76   86   37                                 ______________________________________                                    

EXAMPLE 3

A series of runs was made with Sample 3 jarosite to determine the effectof leaching time in atmospheric leaches and in pressurized leaches attemperatures above the boiling point of the leach solution (at 130° C.and 180° C.). The atmospheric leaches were conducted at 95°-100° C.using the procedure of Example 1. The atmospheric results are given inTable 3A. The pressurized leaches were conducted using Samples 1 and 3in a 500-milliliter reactor fabricated from nickel pipe mounted in ahorizontal configuration. Lifters were installed on the internal surfaceparallel to the axis. The reactor was rotated about its horizontal axisto provide agitation to the reaction mixture. A thermocouple set in athermowell was inserted into the reaction mixture to monitor thetemperature. The contents of the reactor were heated to temperaturewithin about 3 minutes by using external Meker gas burners. The leachmixture was maintained at temperature for the specified time. Thereactor was then quenched in water to lower the temperature rapidly. Thepressurized results are given in Table 3B.

                  TABLE 3A                                                        ______________________________________                                                   Sample 3                                                           Run.sup.(a)                                                                              Leaching Time,                                                                            Extraction, %                                          No.        hours       Ag        Pb  Zn                                       ______________________________________                                        6.sup.(b)  0.5         54        44  21                                       6          1.0         58        56  22                                       6          3.0         61        79  29                                       19.sup.(c) 3.0         57        76  30                                       6          5.0         72        84  28                                       19         5.0         73        86  34                                       19         8           73        82  32                                       19         12          75        84  36                                       19         24          74        85  33                                       ______________________________________                                         .sup.(a) 330 g/l CaCl.sub.2, pH 1.8-3.5                                       .sup.(b) Run 6 had 25 g/l K                                                   .sup.(c) Run 19 had 0 g/l K                                              

                  TABLE 3B                                                        ______________________________________                                        Run      Leaching Time,                                                                            Extraction, %                                            No.      minutes     Ag         Pb   Zn                                       ______________________________________                                        Sample 1 Jarosite (130° C.)                                            20       5           25         9.4  32                                       21       10          53         24   33                                       22.sup.(a)                                                                             10          68         69   42                                       23       20          77         76   46                                       5        30 (autoclave)                                                                            (84)       (80) (45)                                     Sample 3 Jarosite (180° C.).sup.(b)                                    24.sup.(c)                                                                             0           72         57   28                                       25       2.5         69         69   31                                       26       5           72         62   31                                       27       10          69         78   32                                       28       30          71         77   33                                       ______________________________________                                         .sup.(a) 20% solids leach. Other leaches: 16.8% solids.                       .sup.(b) Nickel bomb reactor, 330 g/l CaCl.sub.2, pH 1.8-3.5, 25 g/l K.       .sup.(c) Heat to temperature, cool immediately. Approximately 3 min heat      up.                                                                      

EXAMPLE 4

A series of runs was made using Sample 1 jarosite to determine theeffect of temperature on the amount of solids or pulp density which canbe effectively leached. The leach was conducted using 330 g CaCl₂ /1, at95 to 100° C., for 6 hours, and 1.8 to 3.5 pH. The results are given inTable 4A.

Sample 1 and Sample 3 jarosite materials were leached in the nickel bombreactor of Example 3 with 330 g CaCl₂ /1, pH 1.8-3.5 at temperature for10 minutes. The leach of Sample 1 also contained 25 g/l K. Theseresidues were washed with hot CaCl₂ brine solution (330 g CaCl₂ /1). Theeffect of a single wash is shown in Table 4B. The effect of cumulativewashings on two residues is given in Table 4C.

                  TABLE 4A                                                        ______________________________________                                                   Solubilized, %                                                     Pulp Density Zinc   Silver      Iron Lead                                     ______________________________________                                        30% solids   49     25          1    52                                       30% solids   43     38          1    53                                       20% solids   49     76          1    90                                       15% solids   46     68          1    95                                       10% solids   46     72          1    90                                       ______________________________________                                    

                  TABLE 4B                                                        ______________________________________                                             Initial  Extraction, %                                                   Run  Leach %  Ag            Pb                                                No.  Solids.sup.(a)                                                                         Original Rewash Original                                                                             Rewash Zn                                ______________________________________                                        Sample 1 jarosite (130° C.)                                            22   20       68       83     69     79     42                                29   30       68       84     60     78     37                                30   35       66       82     52     72     38                                31   40       69       82     57     77     40                                32   45       Slurry too thick to agitate                                     33   50       Slurry too thick to agitate                                     Sample 3 jarosite (180° C.)                                            27   16.8     69       --     78     --     32                                34   20       71       --     84     --     33                                35   30       70       74     29     70     30                                36   35       66       72     18     83     31                                37   40       70       76     19     83     31                                38   45       Slurry too thick to agitate                                     39   50       Slurry too thick to agitate                                     ______________________________________                                         .sup.(a) % solids is the weight % of jarosite feed in the initial leach       slurry. Ca(OH).sub.2 added to maintain the leach pH is not included.     

                  TABLE 4C                                                        ______________________________________                                                     Wash           Cumulative                                                     Cumulative     Washing,                                                       Solution       Efficiency, %                                     Filter Cake  Displacements  Ag     Zn                                         ______________________________________                                        Sample 1 residue                                                                           0              0      0                                                       0.6            42.8   47.1                                                    1.2            78.5   83.6                                                    1.9            89.3   91.2                                       Sample 3 residue                                                                           0              0      0                                                       1.6            66.0   60.0                                                    3.25           96.9   94.1                                                    4.9            (100)  98.6                                                    6.6            (100)  99.4                                       ______________________________________                                    

EXAMPLE 5

Closed-cycle process simulation runs were made. Sample 3 feed was usedin Runs 40 and 41 and Sample 1 feed was used in Run 42. The same CaCl₂leach procedure was followed as in Example 1 except the leach time foreach was 30 minutes. The temperature for Run 40 was 180° C. and was 130°C. for Runs 41 and 42. Each leachate from the CaCl₂ leaches wasseparated from solid residue by filtration. In Run 40, a stainless steelautoclave was used which might have resulted in some cementation ofsilver during the leach. The "Cycle No.", e.g. C-1, refers to the numberof times spent brine was recycled to the CaCl₂ leach step. Results aregiven in Table 5A.

The filtrates were cemented with zinc to recover silver and copper. Thecementation was accomplished by bringing the filtrate to temperature ina beaker. Zinc dust was added and the slurry was agitated for one hour.The resulting slurry was filtered and washed with three 50-ml portionsof deionized water. The effect of reaction time and zinc requirement onthe recovery of metals by cementation with zinc powder is shown in Table5B.

                  TABLE 5A                                                        ______________________________________                                        Run   Liquor   Stoichiometric        Barren                                   De-   From     Zn Addition for       Solution,                                scrip-                                                                              Run No.  Ag and Cu (II)                                                                            Cementation, %                                                                          g/l                                      tion  (Cycle)  Cementation Ag  Cu   Pb   Ag   Cu                              ______________________________________                                        A     40       5.9         84  42   --   0.004                                                                              0.042                           B     41, (C-1)                                                                              2.8         60.sup.a                                                                          98   18   0.004                                                                              0.001                           C     41, (C-2)                                                                              14.2        90  66.sup.a                                                                           3    0.001                                                                              0.002                           D     41, (C-3)                                                                              0.7         95  24   0.1  0.001                                                                              0.002                           E     42, (C-1)                                                                              1.2         94  99.4 0.8  0.002                                                                              0.002                           F     42, (C-2)                                                                              1.3         98  99   21   0.001                                                                              0.004                           G     42, (C-3)                                                                              7.3         98  97   99.6.sup.b                                                                         0.001                                                                              0.002                           ______________________________________                                         Other conditions: Temperature 60° C.                                   Time 1 hour                                                                   pH 2.6-5.0                                                                    .sup.a Low value due to low initial concentration in feed to cementation.     .sup.b Anomalous value.                                                  

                  TABLE 5B                                                        ______________________________________                                                      Reaction Cementa-  Barren                                       Stoichiometric Zn                                                                           Time,    tion, %   Solution, g/l                                for Ag & Cu   minutes  Ag      Cu  Ag     Cu                                  ______________________________________                                        1.0           15       62      73  0.008  0.042                                             30       71      70  0.006  0.046                                             60       86      66  0.003  0.051                               1.5           15       67      68  0.007  0.048                                             30       76      68  0.005  0.049                                             60       90      74  0.002  0.040                               2.0           15       71      70  0.006  0.045                                             30       81      77  0.004  0.035                                             60       90      85  0.002  0.023                               ______________________________________                                        Conditions:                                                                           Temperature:                                                                             60° C.                                                      Feed       0.021 g/l Ag, 0.152 g/l Cu,                                        Solution:  3.15 g/l Pb, 6.3 g/l Zn, 0.012 g/l Fe                                         (from autoclave leach of Sample 3,                                            330 g CaCl.sub.2 /l, 180° C. for 1 hour,                               25 g/l K, 25.1 wt. % solids, pH 1.8-                                          3.5).                                                              Zinc       1.0 stoich. = 0.163 g Zn/l                                         Requirement:                                                                             (0.0064 g/l for Ag, 0.157 g/l for Cu).                             Final pH:  5.1-5.2                                                    ______________________________________                                    

EXAMPLE 6

Filtrates from the zinc cementations of Example 5 were each brought totemperature and contacted with hydrogen sulfide gas sufficient toprovide a solution emf below 0.0 millivolts versus a standard calomelelectrode (SCE). This quantity of H₂ S resulted in final solution pH'sof 0.0 to 0.2. After the indicated reaction time, the slurry wasfiltered and the solid was washed with three 50-ml portions of deionizedwater.

                  TABLE 6                                                         ______________________________________                                              From     Run      Solution      Pb                                      Desig-                                                                              Run No.  Time,    emf, mv  Final                                                                              Precipitation                           nation                                                                              (Cycle)  minutes  Initial                                                                             Final                                                                              pH   %                                     ______________________________________                                        A     40       15       +327  -74  1.1  38                                    B     41, (C-1)                                                                              15       +297  -26  0.65 48                                    C     41, (C-2)                                                                              30       +290  -25  0.20 72                                    D     41, (C-3)                                                                              30       +348  -30  --   99.6                                  E     42, (C-1)                                                                              30       +246  +0   --   42                                    F     42, (C-2)                                                                              15       +304  -3   0.28 46                                    G     42, (C-3)                                                                              Data not applicable.                                           ______________________________________                                         Other conditions: Temperature: 60° C.                                  Feed solution pH: 3.5-5.0                                                

EXAMPLE 7

A calcium chloride leach of Sample 3 feed was conducted with a leachsolution of 330 g/l CaCl₂ at an initial solids of 25.1 weight percent ata temperature of about 180° C. (±3° C.) for one hour. The target pH was1.8-3.5 with lb/ton of feed of Ca(OH)₂ added initially to maintain thepH within the target range. The leach mixture was filtered and theleachate solution, maintained at 22° C., was mixed with hydrated lime,Ca(OH)₂, to adjust the pH to the indicated values. The amounts of zincand lead hydroxides precipitated at the particular pH was determined.Results are given in Table 7A.

                  TABLE 7A                                                        ______________________________________                                                  Analysis, g/l                                                                             Precipitation, %                                        pH          Zinc   Lead       Zinc Lead                                       ______________________________________                                        Feed 2.58   8.68   3.31       0.0  0.0                                        3.0         9.33   2.83       0.0  14.5                                       4.0         9.64   2.98       0    10.0                                       5.0         9.58   2.98       0    10.0                                       6.0         10.2   3.28       0    .9                                         7.0         9.7    3.13       0    5.4                                        7.5         5.4    1.4        37.8 57.7                                       8.0         2.06   .125       76.3 96.2                                       9.0         1.89   .055       78.2 98.3                                       9.5         1.47   .441       83.1 86.7                                       10.0        1.47   2.15       83.1 35.0                                       10.7        2.06   2.11       76.3 36.3                                       ______________________________________                                    

The filtrates from Example 6 were each mixed with hydrated lime toadjust the pH to a target final pH of 9.5. The temperature wasmaintained at about 60° C. for 15 minutes. The solid zinc hydroxide wasseparated by filtration using a Buchner funnel. The solids were washedwith three 50-ml portions of deionized water and dried overnight at 100°C. prior to assay. The results are given in Table 7B.

                  TABLE 7B                                                        ______________________________________                                        Test           Ca(OH).sub.2  Zinc     Zinc                                    Desig-                                                                              From     Required Feed Assay, g/l                                                                             Precipitation                           nation                                                                              Test No. g/l      pH   Feed Final %                                     ______________________________________                                        A     40       19.8     0.55 5.39 (1.15).sup.a                                                                        (80)                                  B     41, (C-1)                                                                               7.2     1.0  4.77 1.47  69                                    C     41, (C-2)                                                                              18.0     0.55 5.16 1.00  80                                    D     41 (C-3) --       0.40 4.98 0.88  82                                    E     42, (C-1)                                                                              30.0     0.30 6.24 1.18  81                                    F     42, (C-2)                                                                              --       0.30 7.47 2.21  70                                    G     42, (C-3)                                                                              19.7     1.37 3.35 1.89  62                                    ______________________________________                                         .sup.a Value calculated from solids assays.                              

EXAMPLE 8

Filtrate from Example 7 was mixed with calcium hydroxide and heated atboil. The ammonia evolved was recovered in the distillate or bycondensing or scrubbing the off-gas. The pH was maintained between about8.8 and 10.5 with Ca(OH)₂. Results showing essentially complete NH₃recovery is possible by boiling the solution at the indicated pH areprovided in Table 8. Some of the zinc or lead present in the feedsolution precipitated during the lime boil step and those values aregiven in Table 8 as residuals precipitated.

                                      TABLE 8                                     __________________________________________________________________________    From                  Volume           NH.sub.4                                                                            Residuals                        Run No.    Target                                                                            Ca(OH).sub.2                                                                         Reduction                                                                           NH.sub.4 Assay, g/l                                                                      Volatilized                                                                         Precipitated, %                  Designation                                                                         (Cycle)                                                                            pH  Required, g/l                                                                        %     Feed                                                                             Final                                                                            Distillate                                                                         %     Zn Pb                            __________________________________________________________________________    A     40   10.5                                                                              --     33    1.76                                                                             0.01                                                                             3.02 99.8  48 5                             B     41, (C-1)                                                                          10.5                                                                              3.2    27    1.47                                                                             -- 3.94 100   83 10-12                         C     41, (C-2)                                                                          8.75                                                                              2.4    6.8   -- -- 11.4 (100) .sup.                                                                          4 1                             D     41, (C-3)                                                                          8.75                                                                              --     --    not assayed                                                                              --    -- --                            E     42, (C-1)                                                                          8.75                                                                              0.0    8.8   2.94                                                                             1.06                                                                             21.1 72    48 1                             F     42, (C-2)                                                                          10.5                                                                              1.9    10    not assayed                                                                              --    81 7                             G     42, (C-3)                                                                          8.60                                                                              2.0    4.3   not assayed                                                                              --    74 4                             __________________________________________________________________________

EXAMPLE 9

Ammonia recovery into a high concentration solution from a CaCl₂solution by boiling with Ca(OH)₂ was determined by preparing a syntheticfeed solution. This solution contained 450 g/l CaCl₂ and 50 g/l NH₃ asNH₄ Cl. A volume of solution was placed in a sealed distillationapparatus with provision for incremental sampling of the distillate.Sufficient Ca(OH)₂ was added to liberate 70% of the contained NH₃ andthe distillation was performed with periodic collection and NH₃ analysesof the distillates. The test data are summarized in Table 9. Thedistillation time of 43 minutes produced a 27% volume reduction and a53% recovery of NH₃ in the distillate (75% of the theoretical recoverybased on the Ca(OH)₂ addition). Initial distillate contained over 300g/l NH₃, demonstrating that the lime boil step can recover a highconcentration NH₃ product. The usable NH₃ concentration should be in the200 g/l range as shown by the Distillate No. 1+2+3.

                                      TABLE 9                                     __________________________________________________________________________                                    Distillate Volume,                                                                        NH.sub.3 Evolved,                          Time Volume                                                                             NH.sub.3                                                                         CaCl.sub.2                                                                        NH.sub.3                                                                            % of Feed   % of Theoretical                  Product  minutes                                                                            ml   g/l                                                                              g/l Amount, g                                                                           Increment                                                                           Cumulative                                                                          Increment                                                                           Cumulative                  __________________________________________________________________________    Feed to lime boil                                                                       0   500   (50)                                                                            (450)                                                                             25.0  --    --    0     0                           Distillate No. 1                                                                        7   13.2 305                                                                              --  4.02  2.6   2.6   23.0  23.0                        Distillate No. 2                                                                       14   17.2 173                                                                              --  2.97  3.4   6.1   17.0  40.0                        Distillate No. 3                                                                       20   17.5 146                                                                              --  2.55  3.5   9.6   14.6  54.5                        Distillate No. 4                                                                       29   27.0  70                                                                              --  1.88  5.4   15.0  10.7  65.2                        Distillate No. 5                                                                       30   29.0  37                                                                              --  1.08  5.8   20.8   6.2  71.4                        Distillate No. 6                                                                       43   30.5  20                                                                              --  0.62  6.1   26.9   3.5  75.0                        Distillate total                                                                            134.4       13.12.sup.a                                         Depleted feed 363  -- .sup. (737).sup.b                                       Overall total 497                                                             __________________________________________________________________________     .sup.a Sufficient Ca(OH).sub.2 was added to volatilize 17.5 g NH.sub.3        (100% efficiency); therefore, net efficiency after 43 minutes was 75.0%       (13.12/75 × 100).                                                       .sup.b Calculated value includes CaCl.sub.2 formed from Ca(OH).sub.2 in       lime boil.                                                               

EXAMPLE 10

A zinc ferrite waste was subjected to simultaneous leach-jarositeprecipitation. The ferrite was assayed and found to contain thefollowing components in weight percent: Zn, 13.0; Fe, 33.4; Pb, 0.325;NH₄, 1.12; K, 0.095; and Na, 0.104. Silver was present in the amount of5.06 ounces per ton of waste.

The following procedure was used to leach the solid ferrite waste. Theferrite was mixed with H₂ SO₄ (150 g/l H₂ SO₄) and (NH₄)₂ SO₄ at 20percent weight/weight solids and heated at 90°-95° C. The H₂ SO₄ wasmaintained at the target level indicated in Table 10A by the periodicaddition of H₂ SO₄ solution (150 g/l H₂ SO₄). The mixture was agitatedfor 24 hours with samples taken periodically as indicated in Table 10B.The slurry was filtered to remove solids and the leachate was analyzedwith the results given in Table 10B.

                  TABLE 10A                                                       ______________________________________                                                  Run No.                                                             Variable conditions 43     44    45   46    47                                ______________________________________                                        Leach solution                                                                Initial H.sub.2 SO.sub.4,                                                                 g/l     100    100   100  150   100                                           lb/ton  750    750   750  1125  750                               Initial (NH.sub.4).sub.2 SO.sub.4,                                                        g/l      50    150    84   84    84                                           lb/ton  375    1125  625  625   625                               Target H.sub.2 SO.sub.4,                                                                  g/l      30     30    22   58    30                               ______________________________________                                    

                  TABLE 10B                                                       ______________________________________                                                    Run No.                                                                       43   44       45     46     47                                    Element                                                                              Time, hr   % Extracted                                                 ______________________________________                                        Zn     1          43     28     58   38     33                                       4          57     45     73   58     51                                       8.5        60     53     81   76     59                                       11.5       66     53     82   83     64                                       24         80     65     89   89     75                                Ag     1          2.9    6.2    5.2  5.2    11                                       4          4.4    9      4.4  4.8    4                                        8.5        5.1    6.6    4.7  5.8    4                                        11.5       6.2    5.1    3.3  6      4                                        24         1.8    2      1.8  2.3    1                                 Fe     1          17     17     34   18     18                                       4          6.8    8.8    20   12     8                                        8.5        2.9    3.8    12   16     6                                        11.5       3.5    3      9.2  16     5                                        24         2.7    2.6    5    15     2                                 Pb     1          0      4.4    0    0      4                                        4          1.5    4.7    1.9  0      2                                        8.5        0      0      0.5  0      2                                        11.5       0      3.3    0    0      2                                        24         0      0      0    0      2                                 ______________________________________                                    

EXAMPLE 11

Residues from Runs 45 and 45 of Example 10 were leached with CaCl₂ at95°-100° C. (45a and 46a) and 105° C. (45b and 46b). The procedure ofExample 1 was used to leach the residue from the ferrite leach. Theresults are presented in Tables 11A and 11C.

                  TABLE 11A                                                       ______________________________________                                                           Run No.:                                                                      45a  45b                                                   Element   Time, hr       Extractions, %                                       ______________________________________                                        Zn        1              11     52                                                      3              7.4    36                                                      5              8.3    38                                            Ag        1              22     99                                                      3              19     98                                                      5              24     98                                            Fe        1              0.1    0.5                                                     3              0      0                                                       5              2.2    0.4                                           Pb        1              18     54                                                      3              4.2    50                                                      5              0      35                                            ______________________________________                                    

The recovery for the overall extraction, i.e. the ferrite extraction(given in Table 10B) and the CaCl₂ extraction (given in Table 11A at twotemperatures) for Run 45 are given in Table 11B.

                  TABLE 11B                                                       ______________________________________                                        5 hour Extractions                                                                         Leach(45a)    Leach(45b)                                         Element      Zn     Ag     Fe  Pb  Zn  Ag   Fe   Pb                           ______________________________________                                        Extractions %                                                                 Ferrite Leach                                                                              89     1.8    5   0   89  1.8  5    0                            CaCl.sub.2 Leach                                                                           8.3    24     2.2 0   38  98   0.4  35                           Overall extraction, %                                                                      90     25     7   0   93  98   5    35                           ______________________________________                                    

                  TABLE 11C                                                       ______________________________________                                        Test No.                 46a    46b                                           Element   Time, hr       Extractions, %                                       ______________________________________                                        Zn        1              9.5    3.9                                                     3              23     37                                                      5              23     43                                            Ag        1              8.4    14                                                      3              41     92                                                      5              71     92                                            Fe        1              1.4    0.8                                                     3              1.2    0                                                       5              0.2    0.2                                           Pb        1              18     6.2                                                     3              4.2    54                                                      5              0      39                                            ______________________________________                                    

The recovery for the overall extraction, i.e. the ferrite extraction(given in Table 10B) and the CaCl₂ extraction (given in Table 11C at twotemperatures) for Run 46 are given in Table 11D.

                  TABLE 11D                                                       ______________________________________                                        5 hour Extractions                                                                         Leach(46a)    Leach(46b)                                         Element      Zn     Ag     Fe  Pb  Zn  Ag   Fe   Pb                           ______________________________________                                        Extractions %                                                                 Ferrite Leach                                                                              89     2.3    15  0   89  2.3  15   0                            CaCl.sub.2 Leach                                                                           23     71     0.2 0   43  92   0.2  39                           Overall extraction, %                                                                      92     72     15  0   94  92   15   39                           ______________________________________                                    

EXAMPLE 12

Filtration rate determinations were performed on final slurries from 15of the jarosite leach tests. The filtration data are summarized in Table12. Rates were determined using an Eimco 0.1-ft² vacuum filter leafapparatus, fitted with a medium-weave polypropylene filter cloth. Theapparatus was top-loaded with leach slurry and 18 to 22 inches Hg ofvacuum was applied. The filter cakes were washed first with hot CaCl₂brine followed by water, and the washing rates determined. Theanomalously high wash rates of Runs 48, 49, and 50 probably are due tocake cracking (channeling) and the values are not included in the groupaverages.

The filtration rate averages in Table 12 show that the elevatedtemperature leach slurries, with solids rates of 80 to 230 lbs/ft²/hour, filter two to three times faster than the slurries from 95°-100°C. leaching. Two of the leach slurries were flocculated with Percol 351prior to filtration. Flocculant doses of approximately 200 ppm (solidsbasis) were required to coagulate the slurry solids. The flocculatedfiltration Runs 51 and 52, produced inconsistent results as shown inTable 12.

                                      TABLE 12                                    __________________________________________________________________________    Filtration Rates                                                                                             Cake                                                                  Brine.sup.b                                                                       H.sub.2 O.sup.b                                                                   Thickness                                      Run No.      Solid.sup.a                                                                       Filtrate.sup.b                                                                      Wash                                                                              Wash                                                                              (Inches)                                       __________________________________________________________________________    95-100° C. Leaches                                                     53           9.6 4.0   --  --  1/4                                            54           50  24    7.0 2.1 1/8                                            51           99  30    10.6                                                                              9.0 3/16                                           55           31  9.3   3.2 6.0 1/8                                            52           31  9.5   3.5 3.6 1/4                                             1           44  18    4.6 4.5 5/16                                           50           81  42    (32) .sup.                                                                        (30) .sup.                                                                        3/16                                           Average      56  22    5.8 5.0 3/16                                           (excluding Run 53)                                                            130° C. Autoclave Leaches                                              56           228 95    16  16  5/16                                           48           83  43    31  20  1/8                                            57           91  35    6.8 5.9 1/4                                            58           120 42    13  15  3/16                                           59           130 54    5.8 5.8 1/4                                             5           203 83    7.9 6.7 3/16                                           Average      142 59    9.9 9.9 3/16-1/4                                       180 and 225° C. Leaches                                                49           111 44    21  27  1/4                                            60           75  30    7.3 11  1/4                                            __________________________________________________________________________     .sup.a Solids filtration rate units: lbs/ft.sup.2 /hour.                      .sup.b Liquid filtration rate units: gal/ft.sup.2 /hour.                 

Example 13

Calcium carbonate was used for pH control in a leach of Sample 3 feed.The leach was conducted for 10 minutes at 180° C. in the nickel bombreactor described in Example 3. The run was made under the followingconditions and the results are given in Table 13.

Feed : Sample 3

Leach Solution 330 g/l CaCl₂, Sp.Gr. 1.24,(6.0 N Cl)

K in Solution : 25 g/l initial

Initial % Solids : 20.0% (wt)

Target Leach pH : 1.8-3.5

CaCO₃ required: 283 lb/ton, added initially (Equiv. to 210 lb/tonCa(OH)₂)

Temperature : 180° C. (performed in nickel bomb reactor)

Heat-up Time : 3.5 minutes

Time : 10 minutes

Final Slurry pH : 5.0

                  TABLE 13                                                        ______________________________________                                        Extraction*, %                                                                Zn       Ag            Fe     Pb                                              ______________________________________                                        25.6     53.2          0.2    43.6                                            ______________________________________                                         *metal extractions were calculated from feed and residue assays.         

EXAMPLE 14

Several leaches were conducted at different chloride levels. Leachsolutions of 1.8 normal, 6 normal and 10 normal chloride were used. Theconditions were similar to those of Example 1. The results are given inTable 14.

                  TABLE 14                                                        ______________________________________                                        Leach Solution                                                                g/l CaCl.sub.2                                                                          100 (1.8  .sub.-- NCl.sup.-)                                                              330 (6  .sub.-- NCl.sup.-)                                                               550 (10  .sub.-- NCl.sup.-)                  Feed      Sample 1    Sample 2   Sample 2                                     ______________________________________                                        Extraction, %                                                                 Ag        15          70         74                                           Pb         8          84         89                                           Zn        38          41         42                                           ______________________________________                                    

EXAMPLE 15

A direct recycle of leach filtrate to the next stage of leaching wasconducted to determine the effect of recycle and impurities build-up onthe recovery of values. Five leach cycles were conducted using Sample 3as feed. The first three cycles used the strongly agitated 2-literautoclave, while the final two stages were performed in the less wellagitated nickel bomb reactor. The chloride concentration of the solutionwas determined before and after each cycle and was maintained at 6normal Cl (330 g/l CaCl₂) by addition of CaCl₂, if required. The 16.8%solids leaches were performed for one hour at 180° C. and a pH of 1.8 to4.0 by initial addition of Ca(OH)₂. The leach filtrate from one cyclewas advanced to the next cycle with no intervening solution treatmentsother than reestablishing 6 N Cl concentration, if required.

The results in Table 15 show a steady decrease in the apparent Pbextractions with leach cycling. From cycle 1 to 4, the concentration ofPb in solution increased from 15 to 45 g/l (calculated), whichapproaches Pb saturation in 6 N CaCl₂ brine at 50° to 70° C. Leachslurries were cooled to this temperature range prior to removal from thereactor and filtration. It was found that the reduced Pb extractions incycles two through five were due to saturation of the cooled solutions,causing PbCl₂ to crystallize in the residue solids. The standard residuewashing procedure, two to three cake displacements with 80° C. CaCl₂brine followed by an equal amount of water, was not effective in totallyremoving the PbCl₂ from the residues, thus producing low extractionvalues.

The true leach cycle extractions for Pb ("Pb-corr" column in Table 15)were determined by repulping the residues in 80° to 90° C. CaCl₂ todissolve the residual PbCl₂ prior to reassay of the solids. Thecorrected Pb extractions show that leach cycling does not affect theefficiency of the CaCl₂ leach significantly, provided that the leachslurry is filtered and washed under conditions (temperature and washvolume) which assure complete dissolution of marginally soluble speciessuch as PbCl₂.

                                      TABLE 15                                    __________________________________________________________________________    Run                                                                              Cycle                                                                              Reactor                                                                            Extraction, %                                                                             Filtrate Assay, g/l                                  No.                                                                              No.  Type Zn                                                                              Ag                                                                              Pb Pb-corr.sup.(1)                                                                    Zn  Ag   Pb                                          __________________________________________________________________________    12 1    A    37                                                                              76                                                                              86 86    7.0                                                                              0.071                                                                              (15)                                        12a                                                                              2    A    31                                                                              76                                                                              80 87   11.6                                                                              0.135                                                                              (26)                                        12b                                                                              3    A    35                                                                              75                                                                              76 84   18.9                                                                              0.191                                                                              (36)                                        12c.sup.(2)                                                                      4    B    32                                                                              72                                                                              72 84   (25.4)                                                                            (0.264)                                                                            (45)                                        12d                                                                              5    B    31                                                                              72                                                                              45 81   (21.1)                                                                            (0.253)                                                                            (35)                                        __________________________________________________________________________     Values in parentheses are calculated.                                         .sup.(1) Extractions calculated from rewashed residues. The Pbcorr            (corrected) column presents the actual extraction achieved in the leach.      .sup.(2) Filtrate was diluted to 67% strength prior to advancing to 12d       leach. Dilution occurred when washing slurry from bomb reactor.          

Example 16

A process simulation using closed-cycle steps was performed using Sample1 as feed. The following process steps were included: CaCl₂ leaching,Ag/Cu cementation with zinc, Pb precipitation as sulfide, Znprecipitation as hydroxide, NH₃ evolution by lime boil, and recycle ofthe processed CaCl₂ solution to the next stage of leaching. The values(Ag, Pb, Zn) were removed from the leach solutions prior to recycle tothe next leach. The results are given in Table 16.

                  TABLE 16                                                        ______________________________________                                               Cycle   Element                                                        Process Step                                                                           No.       Zn    Ag   Pb   Cu   Fe    NH.sub.4                        ______________________________________                                                     Extraction, %:                                                   CaCl.sub.2 Leach                                                                       1         46    94   94   47   13 (?)                                         2         48    97   92   51   0.01                                           3         38    96   61.sup.a                                                                           18   0.01                                           Average   44    96   82   39   4                                              Highest   48    97   94   51   0.01                                               Precipitation, %:.sup.b                                          Ag/Cu    1         --    94    0.8 99                                         Cementation                                                                            2         --    98   21   99                                                  3               98   99.6 97                                                  Average         97   41   98                                                  Highest         98   99.6 99                                                      Precipitation, %:.sup.b                                          Lead     1         --    22   42   84                                         Precipitation                                                                          2         --     5   46   53                                                  3                0    0    0                                                  Average         13   44   68                                                  (excluding                                                                    Cycle 3)                                                                      Highest         22   46   84                                                      Precipitation, %:.sup.b                                          Zinc     1         81    --   55   --                                         Precipitation                                                                          2         70    --    0.4 --                                                  3         62    --    9   --                                                  Average   71         22                                                       Highest   81         55                                                           Precipitation, %:                                                                        Evolution, %                                          Lime Boil                                                                              1         48    --    0.6 --         72                                       2         81    --    7   --         --                                       3         74    --    4   --         --                                       Average   68          4              --                                       Highest   81          7              72                              ______________________________________                                         .sup.a Corrected value after rewash of residue. Extraction calculated fro     original residue (incompletely washed) was 12.0%.                             .sup.b % of specie in inlet liquor to particular process step.           

It will be understood that the above description of the presentinvention is susceptible to various modifications, changes andadaptations and the same are intended to be comprehended within themeaning and range of equivalents of the appended claims.

What is claimed is:
 1. In a process for recovering metal values fromwaste containing MFe₃ (SO₄)₂ (OH)₆, where M is a monovalent ion, byleaching said waste with an acidic solution of metal chloride in aclosed system, the improvement comprises leaching said waste with asolution comprising calcium chloride at a temperature above theatmospheric boiling point of the solution and under a pressure of atleast the superatmospheric autogenous pressure which develops as thesystem is heated.
 2. The process of claim 1 wherein said temperature isin excess of about 110° C.
 3. The process of claim 1 wherein saidsolution has a pH of about 1.5 to about 3.5.
 4. The process of claim 1wherein said temperature is between about 120° C. and about 300° C. 5.The process of claim 1 wherein said calcium chloride concentration isbetween about 1.0 molar and the saturation point of the solution.
 6. Theprocess of claim 1 wherein said temperature is between about 150° C. andabout 220° C.
 7. The process of claim 1 wherein potassium is present ata concentration greater than about 0.5 grams per liter of said solution.8. The process of claim 3 wherein said pH is maintained by adding acompound selected from the group consisting of calcium oxide, calciumhydroxide, calcium carbonate and mixtures thereof.
 9. The process ofclaim 1 wherein chloride is provided by adding calcium chloride to saidsolution.
 10. A process for recovering metal values from wastecontaining MFe₃ (SO₄)₂ (OH)₆ where M is a monovalent ion, wherein saidprocess comprises contacting, in a closed system, said waste at atemperature of between about 120° C. and about 200° C. and a pressure ofat least the superatmospheric autogenous pressure which develops as thesystem is heated with a solution containing between about 2.0 and about4.0 molar calcium chloride wherein said solution has a pH of betweenabout 1.5 and about 3.5 which is maintained by the addition of a calciumcompound selected from the group consisting of calcium oxide, calciumhydroxide, calcium carbonate and mixtures thereof.
 11. The process ofclaim 10 wherein said waste comprises ammonium jarosie formed bysubjecting a material containing zinc ferrite to a leach said processconsisting essentially of:(a) mixing an aqueous slurry of said materialwith sulfuric acid and a source of ammonium ions to form a sulfuric acidleach mixture; (b) heating said sulfuric acid leach mixture to form asolid containing ammonium jarosite and a liquid containing zinc sulfate;and (c) conveying said solid into contact with said solution of calciumchloride.
 12. The process of claim 10 wherein said leaching provides aliquid leachate and a solid residue and wherein said process furthercomprises:(a) separating said liquid leachate from said solid residuewherein said solid residue comprises an iron oxide; (b) contacting saidliquid leachate with a reducing metal to reduce silver cations containedin said leachate to metallic silver and then recovering said metallicsilver from the liquid phase; (c) recovering zinc from the liquid phaseof step (b) using a zinc recovery process to provide a liquid solutionsubstantially free of zinc; and (d) adjusting the pH of the liquidsolution from step (c) to above about 9 by adding a basic material andthen heating the solution to provide a vapor containing ammonia.
 13. Theprocess of claim 12 wherein said zinc is recovered in step (c) by aprocess comprising extracting said zinc by contacting the liquid phasecontaining said zinc with an extractant to remove said zinc from theliquid phase to a second phase.
 14. The process of claim 12 wherein thebasic material from step (d) is selected from the group consisting ofcalcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof.15. The process of claim 14 wherein said basic material consistsessentially of calcium oxide.
 16. The process of claim 12 wherein saidreducing metal is selected from the group consisting of lead and zinc.17. The process of claim 16 wherein said reducing metal is lead andwherein the process further comprises the steps (i) mixing the liquidphase remaining in step (b) after removing said metallic silver with asulfide compound to precipitate the lead as lead sulfide, and (ii)removing substantially all of said lead sulfide from the solution andthen conveying said substantially lead-free solution to the zincrecovery process.
 18. The process of claim 16 wherein said reducingmetal is lead and wherein the process further comprises the steps of:(a)contacting the substantially zinc-free liquid phase from step (c) withsufficient calcium oxide, calcium hydroxide, calcium carbonate ormixtures thereof to provide a solution pH of between about 8 and about 9to precipitate lead hydroxide; (b) removing the precipitated leadhydroxide from the liquid phase; (c) adding sufficient calcium oxide,calcium hydroxide, calcium carbonate or mixtures thereof to the liquidphase from step (ii) to increase the pH to above about 10; and (d)increasing the temperature of the resulting solution to about theboiling point of said solution.
 19. The process of claim 12 wherein saidzinc is recovered by contacting said liquid phase from step (b) withcalcium oxide, calcium hydroxide, calcium carbonate or mixtures thereofto precipitate said zinc as zinc hydroxide and separating said zinchydroxide precipitate from the liquid phase.
 20. A process forrecovering metal values from waste containing MFe₃ (SO₄)₂ (OH)₆, where Mis a monovalent ion, said process comprising:(a) mixing an aqueousslurry containing a material comprising zinc ferrite, sulfuric acid anda source of ammonium ions to form a sulfuric acid leach mixture; (b)heating said leach mixture to form a solid phase containing ammoniumjarosite and a liquid phase containing zinc sulfate; (c) separating saidsolid phase and said liquid phase; (d) contacting said separated solidphase with a solution comprising between 1.0 and 5.0 molar calciumchloride in a closed system at a temperature between about 110° C. andabout 300° C. and a pressure of at least the superatmospheric autogenouspressure which develops as the system is heated wherein said solutionhas a pH of between about 1.5 and about 3.5 to form a liquid leachateand a solid residue; (e) separating said liquid leachate from said solidresidue; (f) contacting said liquid leachate with a reducing metalselected from the group consisting of lead and zinc to reduce silvercations contained in said leachate to metallic silver; (g) separatingsaid metallic silver from the liquid phase as a cement; (h) recoveringzinc contained in said liquid phase by contacting said liquid phase withan extractant which selectively removes said zinc from said liquid phaseto a second phase; (i) contacting the substantially zinc-free liquidphase from step (h) with sufficient calcium oxide, calcium hydroxide,calcium carbonate or mixtures thereof to provide a solution pH ofbetween about 8 and 9 to precipitate lead hydroxide; (j) removing theprecipitated lead hydroxide from the liquid phase; (k) adding sufficientcalcium oxide, calcium hydroxide, calcium carbonate or mixtures thereofto the liquid phase from (j) to increase the pH to above about 10; (l)increasing the temperature of the resulting solution to form a vaporcontaining ammonia and a liquid residue containing calcium chloridebrine; (m) condensing the vapor from step (1) to provide a solutioncontaining ammonium hydroxide; and (n) recycling the liquid residue fromstep (1) to the calcium chloride leaching of step (d).